Method for recovering metal values from concentrates of sulfide minerals

ABSTRACT

A method of recovering precious metal values from refractory sulfide ores is provided. The method includes the steps of separating clays and fines from a crushed refractory sulfide ore, forming a heap from the refractory sulfide ore, producing a concentrate of refractory sulfide minerals from the separated fines and adding the concentrate to the heap, bioleaching the heap to thereby oxidize iron sulfides contained therein, and hydrometallurgically treating the bioleached ore to recover precious metalvalues contained therein.

[0001] This is a continuation of copending application Ser. No.09/709,765 filed Nov. 10, 2000, which is a continuation of copendingapplication Ser. No. 08/950,279, filed Oct. 14, 1997, now U.S. Pat. No.6,146,444, which is a continuation of copending application Ser. No.08/476,444, filed on Jun. 7, 1995, now U.S. Pat. No. 5,676,733, which isa continuation in part of copending application Ser. No. 08/343,888,filed on Nov. 16, 1994, now U.S. Pat. No. 5,573,575, which is acontinuation in part of copending application Ser. No. 08/161,742, filedon Dec. 3, 1993, now U.S. Pat. No. 5,431,717.

BACKGROUND OF THE INVENTION

[0002] 1. Field of the Invention

[0003] The present invention relates to the recovery of metal valuesfrom refractory sulfide and refractory carbonaceous sulfide ores.

[0004] 2. Description of the Prior Art

[0005] Gold is one of the rarest metals on earth. Gold ores can becategorized into two types: free milling and refractory. Free millingores are those that can be processed by simple gravity techniques ordirect cyanidation. Refractory ores, on the other hand, are not amenableto conventional cyanidation treatment. Such ores are often refractorybecause of their excessive content of metallic sulfides (e.g., pyrite)and/or organic carbonaceous matter.

[0006] A large number of refractory ores consist of ores with a preciousmetal such as gold occluded in iron sulfide particles. The iron sulfideparticles consist principally of pyrite and arsenopyrite. Precious metalvalues are frequently occluded within the sulfide mineral. For example,gold often occurs as finely disseminated sub-microscopic particleswithin a refractory sulfide host of pyrite or arsenopyrite. If the goldremains occluded within the sulfide host, even after grinding, then thesulfides must be oxidized to liberate the encapsulated precious metalvalues and make them amenable to a leaching agent (or lixiviant).

[0007] A number of processes for oxidizing the sulfide minerals toliberate the precious metal values are well known in the art. One knownmethod of oxidizing the metal sulfides in the ore is to use bacteria,such as Thiobacillus ferrooxidans, Sulfolobus, Acidianus species andfacultative-thermophilic bacteria in a microbial pretreatment. Theforegoing microorganisms oxidize the iron sulfide particles to cause thesolubilization of iron as ferric iron, and sulfide, as sulfate ion.

[0008] If the refractory ore being processed is a carbonaceous sulfideore, then additional process steps may be required following microbialpretreatment to prevent preg-robbing of the aurocyanide complex or otherprecious metal-lixiviant complexes by the native carbonaceous matterupon treatment with a lixiviant.

[0009] As used herein, sulfide ore or refractory sulfide ore will beunderstood to also encompass refractory carbonaceous sulfide ores.

[0010] A known method of bioleaching carbonaceous sulfide ores isdisclosed in U.S. Pat. No. 4,729,788, issued Mar. 8, 1988, which ishereby incorporated by reference. According to the disclosed process,thermophilic bacteria, such as Sulfolobus and facultative-thermophilicbacteria, are used to oxidize the sulfide constituents of the ore. Thebioleached ore is then treated with a blanking agent to inhibit thepreg-robbing propensity of the carbonaceous component of the ore. Theprecious metals are then extracted from the ore using a conventionallixiviant of cyanide or thiourea.

[0011] Another known method of bioleaching carbonaceous sulfide ores isdisclosed in U.S. Pat. No. 5,127,942, issued Jul. 7, 1992, which ishereby incorporated by reference. According to this method, the ore issubjected to an oxidative bioleach to oxidize the sulfide component ofthe ore and liberate the precious metal values. The ore is theninoculated with a bacterial consortium in the presence of nutrientstherefor to promote the growth of the bacterial consortium, thebacterial consortium being characterized by the property of deactivatingthe preg-robbing propensity of the carbonaceous matter in the ore. Inother words, the bacterial consortium functions as a biological blankingagent. Following treatment with the microbial consortium capable ofdeactivating the precious-metal-adsorbing carbon, the ore is thenleached with an appropriate lixiviant to cause the dissolution of theprecious metal in the ore.

[0012] Problems exist, however, with employing bioleaching processes ina heap leaching environment. These include nutrient access, air access,and carbon dioxide access for making the process more efficient and thusan attractive treatment option. Moreover, for biooxidation, theinduction times concerning biooxidants, the growth cycles, viability ofthe bacteria and the like are important considerations because thevariables such as accessibility, particle size, settling, compaction andthe like are economically irreversible once a heap has been constructed.As a result, heaps cannot be repaired once formed, except on a limitedbasis.

[0013] Ores that have a high clay and/or fines content are especiallyproblematic when processing in a heap leaching or heap biooxidationprocess. The reason for this is that the clay and/or fines can migratethrough the heap and plug channels of air and liquid flow, resulting inpuddling; channelling; nutrient-, carbon dioxide-, or oxygen-starving;uneven biooxidant distribution, and the like. As a result, large areasof the heap may be blinded off and ineffectively leached. This is acommon problem in cyanide leaching and has lead to processes of particleagglomeration with cement for high pH cyanide leaching and with polymersfor low pH bioleaching. Polymer agglomerate aids may also be used inhigh pH environments, which are customarily used for leaching theprecious metals, following oxidative bioleaching of the iron sulfides inthe ore.

[0014] Biooxidation of refractory sulfide ores is especially sensitiveto blocked percolation channels by loose clay and fine material becausethe bacteria need large amounts of air or oxygen to grow and biooxidizethe iron sulfide particles in the ore. Air flow is also important todissipate heat generated by the exothermic biooxidation reaction,because excessive heat can kill the growing bacteria in a large, poorlyventilated heap.

[0015] Ores that are low in sulfide or pyrite, or ores that are high inacid consuming materials such as calcium carbonate or other carbonates,may also be problematic when processing in a heap biooxidation. Thereason for this is that the acid generated by these low pyrite ores isinsufficient to maintain a low pH and high iron concentrate needed forbacteria growth.

[0016] A need exists, therefore, for a heap bioleaching technique thatcan be used to biooxidize precious metal bearing refractory sulfide oresand which provides improved air and fluid flow within the heap. Inaddition, a need exists for a heap bioleaching process in which oresthat are low in sulfide minerals, or ores that are high in acidconsuming materials such as calcium carbonate, may be processed.

[0017] A need also exists for a biooxidation process which can be usedto liberate occluded precious metals in concentrates of refractorysulfide minerals. Mill processes that can be used for oxidizing suchconcentrates include bioleaching in a stirred bioreactor, pressureoxidation in an autoclave, and roasting. These mill processes oxidizethe sulfide minerals in the concentrate relatively quickly, therebyliberating the entrapped precious metals. However, unless theconcentrate has a high concentration of gold, it does not economicallyjustify the capital expense or high operating costs associated withthese processes. And, while a mill bioleaching process is the leastexpensive mill process in terms of both the initial capital costs andits operating costs, it still does not justify processing concentrateshaving less than about 0.5 oz. of gold per ton of concentrate, whichtypically requires an ore having a concentration greater than about 0.07oz. of gold per ton. Therefore, a need also exists for a process thatcan be used to biooxidize concentrates of precious metal bearingrefractory sulfide minerals at a rate comparable to a stirred tankbioreactor, but that has capital and operating costs more comparable tothat of a heap bioleaching process.

[0018] In addition to concentrates of precious metal bearing sulfideminerals, there are many sulfide ores that contain metal sulfideminerals that can potentially be treated using a biooxidation process.For example, many copper ores contain copper sulfide minerals. Otherexamples include zinc ores, nickel ores, and uranium ores. Biooxidationcould be used to cause the dissolution of metal values such as copper,zinc, nickel and uranium from concentrates of these ores. The dissolvedmetal values could then be recovered using known solvent extractiontechniques, iron cementation, and precipitation. However, due to thesheer volume of the sulfide concentrate formed from sulfide ores, astirred bioreactor would be prohibitively expensive, and standard heapoperations would simply take too long to make it economically feasibleto recover the desired metal values. A need also exists, therefore, foran economical process for biooxidizing concentrates of metal sulfideminerals produced from sulfide ores to thereby cause the dissolution ofthe metal values so that they may be subsequently recovered from thebioleachate solution.

SUMMARY OF INVENTION

[0019] It is an object of one aspect of the present invention to providea heap bioleaching process of the type described above, wherein therefractory sulfide ore is rendered more susceptible to biooxidation,thereby providing improved recovery of the precious metal valuescontained within the ore. The method of the present invention achievesthis object by removing the clays and/or fines from the refractorysulfide ore after it is crushed to a size appropriate for a heapleaching process. The heap may then be formed without concern of the airand liquid flow channels in the heap becoming plugged. Further, if theseparated clay and/or fine material has a sufficiently high preciousmetal content, it may be separately treated to recover the preciousmetal values contained therein.

[0020] In another aspect of the present invention, a process forrecovering precious metal values from concentrates of precious metalbearing refractory sulfide minerals is provided. The process comprisesthe steps of (a) distributing a concentrate of refractory sulfideminerals on top of a heap of support material; (b.) biooxidizing theconcentrate of refractory sulfide minerals; (c.) leaching precious metalvalues from the biooxidized refractory sulfide minerals with alixiviant; and (d.) recovering precious metal values from the lixiviant.An advantage of this process is that the rate at which the sulfideminerals biooxidize is much higher than would be observed in atraditional heap bioleaching operation. Despite this high rate ofbiooxidation, however, the initial capital costs and operating costs forthe disclosed process are lower than that associated with a mill typebiooxidation process.

[0021] Gold is the preferred precious metal recovered using the processaccording to the present aspect of the invention. However, otherprecious metals can also be recovered, including silver and platinum.The support material is preferably selected from the group consisting oflava rock, gravel, and coarsely ground ore. Lava rock is particularlypreferred due to its high surface area. As those skilled in the art willimmediately recognize, a number of lixiviants can be used in conjunctionwith the present process, however, thiourea and cyanide are thepreferred, cyanide being a particularly preferred lixiviant.

[0022] In another aspect of the present invention a process is providedfor recovering metal values from sulfide ores. Such ores include, by wayof example, chalcopyrite, sphalorite, nickel sulfide ores, and uraniumsulfide ores. The process according to this aspect of the inventioncomprises (a.) forming a concentrate of metal sulfide minerals; (b.)spreading the concentrate on top of a heap of support material; (c.)biooxidizing the concentrate; and (d.) recovering metal values from thesolution used to biooxidize the metal sulfide minerals. Due to the factthat this process uses a heap of support material for the bioreactor,its capital and operating costs are less than that of a mill bioleachingoperation. However, due to the good air flow in the heap, thebiooxidation rate of the sulfide minerals is quite high and can approachthat of what would be observed in a mill type operation.

[0023] Depending on the sulfide ore from which the concentrate isobtained, the metal values that can be recovered from the processaccording to the present aspect of the invention include copper, zinc,nickel and uranium. The support material used in the present process ispreferably selected from the group consisting of lava rock, gravel, andcoarsely ground rock. Lava rock is particularly preferred due to itshigh surface area.

[0024] The above and other objects, features and advantages will becomeapparent to those skilled in the art from the following description ofthe preferred embodiments.

BRIEF DESCRIPTION OF THE DRAWING

[0025]FIG. 1 is a schematic of a process flow sheet according to apreferred embodiment of the present invention;

[0026]FIG. 2 is a graph illustrating the percent iron leached over timefor various size fractions of ore;

[0027]FIG. 3 is a graph illustrating the percent iron leached per day asa function of time for various size fractions of ore;

[0028]FIG. 4 is a graph illustrating the percent gold recovered from apyrite concentrate milled to −200 mesh as a function of its percentbiooxidation;

[0029]FIG. 5 is a graph illustrating the change in Eh of a column of+{fraction (1/4)} inch ore as a function of time;

[0030]FIG. 6 is a graph illustrating the change in pH as a function oftime for a column of +{fraction (1/4)} inch ore;

[0031]FIG. 7 is a graph illustrating the change in iron concentration inthe effluent of a column of +{fraction (1/4)} inch ore as a function oftime; and

[0032]FIG. 8 is a graph illustrating the biooxidation rate for aconcentrate on lava rock process according to the present inventionversus a stirred tank type process.

DESCRIPTION OF THE PREFERRED EMBODIMENTS

[0033] According to one aspect of the present invention, refractorysulfide ores can be rendered more susceptible to biooxidation in a heapleaching process. This is accomplished by separating the clay and/orfine materials from the refractory sulfide ore after it has been crushedto a size appropriate for heap leaching. In the present embodiment themethod of removal is wet size screening. It will be readily apparent tothose skilled in the art, however, that any other method for separatingthe clay and/or fine material from the refractory ore may be used. Forexample, dry screening and cyclone classifying are well known to thoseskilled in the art.

[0034] By removing the fines and clays from the refractory sulfide ore,the air and liquid flow through the heap is improved. This will reducethe time required to sufficiently biooxidize the iron sulfide particlesin the ore to liberate the precious metal values and make them amenableto subsequent lixiviation with cyanide or thiourea, preferably cyanide.In addition to faster biooxidation, in a well ventilated heap, havinggood fluid flow, it becomes more feasible to change the pH from anacidic pH of 1.0 to 2.0 that is best for biooxidation to a basic pH of10.0 or more needed for cyanide leaching without remaking or restackingthe heap.

[0035] The refractory sulfide ore is preferably crushed to a targetmaximum size in the range of approximately {fraction (1/4)} to 1 inch.Appropriate target maximum particle sizes include ¼, ⅜, ½, and 1 inch.If the ore will pass any of these target particle sizes, it should beamenable to heap leaching. The smaller the particle size, however, thegreater the surface area of the sulfide particles in the ore and, ofcourse, the faster the sulfide particles will be biooxidized. Increasedrecovery of the precious metal values should also result. This, however,must be weighed against the additional cost of crushing the ore to asmaller particle size. The additional amount of precious metal recoveredmay not justify the added cost.

[0036] Of course if the refractory sulfide ore body being treated isalready an appropriate size for heap leaching, no additional crushing isrequired.

[0037] Fines are naturally produced during the crushing process. Thesize of the fines and clays removed from the crushed ore should be aboutminus 60 mesh as a minimum upper limit to about minus {fraction (1/8)}inch as a maximum upper limit. After the clay and fines are separatedfrom the bulk of the ore, a heap is formed with the ore. The heap maythen be treated with a standard bioleaching process to oxidize the ironsulfide particles in the ore and liberate the occluded precious metalvalues, which are preferably gold. Because the majority of the clay andfine materials have been removed, obstruction of the air and liquid flowchannels by these materials is no longer a concern, thereby improvingpercolation leaching of the ore.

[0038] After biooxidation, the precious metal in the pretreated ore canbe extracted using a conventional lixiviant such as cyanide or thiourea,preferably cyanide. Of course, however, as a person of ordinary skill inthe art would recognize, if the refractory sulfide ore is alsorefractory due to carbonaceous matter contained in the ore, additionalprocessing steps must be employed to reduce the preg-robbing propensityof the ore prior to lixiviation. A number of such processes are wellknown in the art.

[0039] For example, the methods used in U.S. Pat. No. 4,729,788 and U.S.Pat. No. 5,127,942, both of which have already been incorporated hereinby reference, can be used. Further, the microbial process for treatingcarbonaceous ores disclosed in U.S. Pat. No. 5,162,105, issued Nov. 10,1992, hereby incorporated by reference, can also be used.

[0040] The fine material that has been separated may contain largeamounts of precious metal values. Indeed the economic value of thesemetal values may be sufficiently high to justify further processing ofthese materials to recover the additional metal values. In aparticularly preferred embodiment of the present invention, theseparated fine material is further processed to recover at least aportion of the precious metal values contained therein.

[0041] To recover the precious metal values from the fine material, thefine material is preferably treated in a mill process to remove the ironsulfide particles from the clay and sand particles. The reason for thisis that, as discussed above, precious metal values, especially gold,often occur as finely disseminated microscopic particles within the ironsulfide particles. These fine sulfide particles, therefore, frequentlycontain a significant portion of the overall precious metal values.Further, because a relatively high percentage of the precious metalvalues in the ore are associated with this fraction of the ore, they canbe economically treated in a mill process.

[0042] As will be recognized by those skilled in the art, a variety ofmethods can be used to separate the iron sulfide particles from theremainder of the fines. These methods include, by way of example only,gravity separation and flotation. If desired, the iron sulfide particlescan be subjected to additional grinding before flotation. Gravityseparation techniques that can be used include shaker tables,hydrocyclones, and spiral classifiers.

[0043] The iron sulfide concentrate, if refractory, is preferablybioleached with bacteria in a tank or mill process to liberate theoccluded precious metal values. Alternatively, the sulfide concentratecan be added back to the heap to allow for a slower heap biooxidationprocess. However, because these particles are typically larger and morehydrophobic than clay particles, they tend to stick more readily to thelarger particles in the heap, and, thus, the problem of obstructedpercolation channels is not encountered. The iron sulfide concentratecan also be treated by a variety of other methods well known in the artsuch as roasting, pressure oxidation, and chemical oxidation. Becausethe concentration of gold or other precious metal values is relativelyhigh in this ore fraction and its overall volume small, all of thesemill processes may be economically utilized.

[0044] If the iron sulfide concentrate is only partially refractory,then it can be directly leached with a lixiviant such as cyanide toremove the nonrefractory gold. The tail from this leaching process couldthen be washed free of cyanide and added to the heap for biooxidation torelease the remaining refractory gold or other precious metal values.

[0045] The fine material removed from the refractory sulfide ore by sizeseparation, and which has also had the iron sulfide particles removedfrom it, may still contain economic values of gold or other preciousmetals. Further, this fine material is likely to be less refractory thanother iron sulfide material if the size has lead to oxidation.Therefore, agglomeration of this material with cement, or otheragglomeration aids that can be used at a high pH, may provide goodrecoveries if leached with cyanide directly.

[0046] The fine material may have sufficient gold value in the case ofhigh grade ore to merit a mill leaching process such as carbon-in-pulpor counter current decantation.

[0047] A more recently preferred embodiment of the present invention isnow described in connection with the process flow sheet illustrated inFIG. 1.

[0048] As can be seen from referring to FIG. 1, a precious metal bearingrefractory sulfide ore is preferably crushed to a target maximum size inthe range of approximately {fraction (1/4)} to 1 inch at crushingstation 10. Preferably the ore is crushed to a target maximum particlesize of ¼, ⅜, ½, or ¾ inch. Of course, if the refractory sulfide orebody being treated is already of an appropriate size for heap leaching,no additional crushing is required.

[0049] As in the present embodiment, the precious metal to be recoveredfrom the ore is typically gold. However, as those skilled in the artwill readily recognize, the method according to the present invention isequally applicable to the recovery of other precious metals, includingsilver and platinum from refractory sulfide ores.

[0050] After the gold bearing refractory sulfide ore is crushed to theappropriate size, the fines in the ore are separated from the crushedore at separation station 12. Preferably the fines are separated using awet or dry screening process. To ensure good air and liquid flow in theheap, fines smaller than about 10 to 30 mesh (Tyler mesh series) shouldbe separated out at separation station 12. The coarse fraction of theore 14, that is the ore greater than about 10 to 30 mesh, will typicallycontain approximately 50% or more of the gold values in the entire oreand comprise about 50 to 75% of the weight of the ore. The fines 16 thathave been separated out will typically contain approximately 30 to 50%of the gold values and comprise approximately 25 to 50% of the weight ofthe initial ore.

[0051] Because of the significant gold values typically contained in thefines 16, the fines are further processed to recover at least a portionof the precious metal values contained therein. This is preferablyaccomplished by producing a concentrate 20 of refractory pyrite mineralsfrom the fines 16 in the pyrite concentration circuit 22. Pyriteconcentrate 20 will typically comprise about 5 to 10% of the initialweight of the ore and about 15 to 30% of its gold values.

[0052] If the ore contains refractory arsenopyrite minerals, thenrefractory pyrite concentrate 20 will also contain these minerals.

[0053] Because, as a general rule, the pyrite particles in the pyriteconcentrate 20 are larger and more hydrophobic than the clay particlesfound in fines 16, the pyrite concentrate 20 can be combined with thecoarse fraction of the ore 14 during heap construction withoutsignificantly impeding fluid and air flow within the heap duringbioleaching. This is because the pyrite particles in pyrite concentrate20 will tend to stick to the larger particles in the coarse fraction ofthe ore, rather than migrating through the heap and causing blocked flowchannels. Pyrite concentrate 20, may also be added to the top of theheap before or after the biooxidation process is already in progress.

[0054] The bacterial oxidation of pyrite generates ferrous sulfate andsulfuric acid in the net reaction summarized by Equation (1). This netreaction can be broken into two distinct reactions, Equations (2) and(3), where Equation (2) is the aerobic reaction catalyzed by bacterialactivity and Equation (3) is the anaerobic reaction occurring at thesurface of the sulfide mineral. Equation (4) is a similar anaerobicreaction occurring at the surface of arsenopyrite minerals.

FeS₂+7/2O₂+H₂O=FeSO₄+H₂SO₄  (1)

14FeSO₄+7H₂SO₄+7/2O₂=7H₂O+7Fe₂(SO₄)₃  (2)

7Fe₂(SO₄)₃+FeS₂+8H₂O=15FeSO₄+8H₂SO₄  (3)

13Fe₂(SO₄)₃+2FeAsS+16H₂O=20FeSO₄+2H₃AsO₄+13H₂SO₄  (4)

[0055] An advantage of adding pyrite concentrate 20 to heap 18 is thatthis fine milled pyrite is more readily oxidized than the pyrite mineralparticles found in. coarse ore 14; thus, the acid produced from theoxidation of the pyrite concentrate can be used to help lower the pH ofthe coarse ore in the heap more quickly. This is especially valuablewhen dealing with ores that are high in acid consuming materials such ascalcium carbonate or other carbonates. Further, by adding the pyriteconcentrate to the top of heap 18, ferric ions produced during itsbiooxidation will migrate to the lower part of the heap where bacterialgrowth may be inhibited due to toxins, which have not been washed fromthe ore early in the biooxidation process, or due to the lack of oxygen.As a result, biooxidation of the pyrite minerals in the lower part ofthe heap will proceed even if bacterial growth is not favored in thisregion.

[0056] There is also an advantage to adding pyrite concentrate 20 to aheap 18 that has been undergoing biooxidation for a long period of time.In the later stages of biooxidation most of the exposed and reactivesulfides will have already been oxidized, resulting in a slow down inthe rate of biooxidation. This slow down in the rate of biooxidation canlead to a drop in iron levels and an increase in pH within heap 18.Addition of a reactive sulfide concentrate can restart an activebiooxidation process that can increase indirect chemical leaching ofimbedded sulfide minerals due to the high ferric levels produced fromthe biooxidation of the sulfide concentrate.

[0057] The preferred methods of producing pyrite concentrate 20 areexplained in detail below in connection with pyrite concentrationcircuit 22.

[0058] After heap 18 is constructed, it may be pretreated using astandard heap biooxidation process to oxidize the iron sulfide particlesin the ore and liberate the occluded precious metal values. And, becausethe majority of the clay and fine materials have been removed,obstruction of the air and liquid flow channels by these materials issignificantly reduced, resulting in improved percolation leaching of theore.

[0059] If the bioleachate solution is recycled during the biooxidationprocess, the biooxidation rate can be improved by using the method ofsolution management disclosed in the U.S. patent application Ser. No.08/329,002, entitled “Method For Improving The Heap Biooxidation Rate OfRefractory Sulfide Ore Particles That Are Biooxidized Using RecycledBioleachate Solution,” which was filed Oct. 25, 1994, by William J.Kohr, Chris Johannson, John Shield, and Vandy Shrader, the text of whichis incorporated herein by reference as if fully set forth.

[0060] Referring again to FIG. 1, pyrite concentration circuit 22 is nowdescribed. Three preferred methods of producing pyrite concentrate 20are illustrated within pyrite concentration circuit 22. These methodsmay be used in combination or in the alternative.

[0061] The fines 16 will typically comprise very fine clay particles,which are typically less than 5 to 20 μm; sand particles; and refractorysulfide particles. The clay particles are very small and veryhydrophilic in comparison to the sand and refractory sulfide particles,making them particularly deleterious to heap bioleaching processes,because they tend to migrate through the heap and plug flow channels asthey swell from the absorption of water. The clay particles are,therefore, preferably removed from the fines 16 so that a concentrate ofrefractory sulfide particles can be produced that can be safely added toheap 18 with minimal obstruction of the flow channels in the heap. Thus,as illustrated in FIG. 1, the first step in each of the preferredmethods of producing pyrite concentrate 20 is the removal of the clayparticles from the fines using clay removal system 24, which ispreferably a hydrocyclone or a settling tank. Of course, however, if theore is a low clay bearing ore, this step may be omitted.

[0062] The set point for the maximum size particle removed in clayremoval system 24 will depend on the distribution of clay particle sizeswithin fines 16. If the set point for the clay removal system is set atless than about 10 μm, a settling tank is the preferred removal methodof separation because hydrocyclones cannot currently make efficientseparations between particle sizes of less than about 10 μm.

[0063] In a high clay ore, clay material 26 separated from the fines 16will typically comprise about 10% of the initial weight of the ore andabout 5 to 10% of its gold values. Further, because of its lowrefractory nature, clay material 26 may be further processed to recoverthe gold values it contains using a traditional cyanide mill leachingprocess such as counter current decantation or carbon-in-pulp. Beforeprocessing clay material 26 in one of these traditional cyanide millleaching processes, however, the pulp density of the clay materialshould be increased using a thickener 28 until a pulp density of about30 to 40% is achieved.

[0064] After the clays have been removed from the fines 16, therefractory pyrite particles are also separated out to form refractorypyrite concentrate 20, which can be added to heap 18 as explained above.The refractory pyrite particles are preferably separated from claydepleted fines 30 using flotation or a gravity separation technique.

[0065] Three preferred methods for separating the refractory sulfideparticles from the clay depleted fines 30 are now described. The firstmethod entails fine grinding the clay depleted fines 16 until a particlesize of less than about −200 mesh is achieved. This is preferablyaccomplished in ball mill 34. The refractory pyrite materials are thenremoved from the material 30 using a flotation cell 36 with a xanthatecollector. The floated pyrite material from flotation cell 36 forms thepyrite concentrate 20.

[0066] A second method of producing pyrite concentrate 20 from material30 comprises separating material 30 into two fractions using ahydrocyclone 38: the first, comprising −200 mesh material 40, and thesecond comprising coarse sand particles, which are greater than about200 mesh, and heavier pyrite particles. The material which is less than200 mesh is further treated in xanthate flotation cell 36 to removerefractory sulfides. The floated refractory sulfides and the coarse sandparticles and heavier pyrite are then recombined to form pyriteconcentrate 20. This method differs from the first pyrite concentrationmethod in that instead of crushing all of material 30 to less than −200mesh, the sand particles greater than 200 mesh and the heavier pyriteminerals in material 30 are simply separated from material 30 and thenadded to the floated pyrite from the −200 mesh material 40.

[0067] The third method of producing pyrite concentrate 20 from claydepleted fines 30 comprises using a gravity technique such as a spiralclassifier or shaker table to remove the heavier sulfide minerals fromthe remainder of material 30.

[0068] The tail material 32, which remains after the refractory sulfidefraction has been removed from the clay depleted fines material 30,comprises approximately 20 to 30% of the initial weight of the ore andabout 5 to 10% of its gold, approximately 85% of which is recoverable ina traditional cyanide mill leaching process such as counter currentdecantation or carbon-in-pulp. Thus, tail material 32 is not veryrefractory and may be treated with clay material 26 in a traditionalmill cyanide leaching process to help improve the overall recovery ofthe process.

[0069] After heap 18 is biooxidized, the precious metal in thepretreated ore can be extracted using a conventional lixiviant such ascyanide or thiourea, preferably cyanide. Of course, however, as a personof ordinary skill in the art would recognize, if the refractory sulfideore is also refractory due to carbonaceous matter contained in the ore,additional processing steps must be employed to reduce the preg-robbingpropensity of the ore prior to lixiviation as explained above.

EXAMPLE 1

[0070] A sample of 16 kg of refractory sulfide ore with approximately0.04 oz/ton of gold and 3.5% of sulfide sulphur was crushed to−{fraction (1/4)} inch. The ore sample was then separated by wetscreening into a +{fraction (1/8)} to −{fraction (1/4)} inch, a +30 meshto −{fraction (1/8)} inch, and a −30 mesh material fractions. The −30mesh material was further separated into a pyrite fraction, a sandfraction, and a clay fraction by gravity separation. The sand fractionwas further processed by fine grinding in a ball mill for about onehour. This material was then floated with xanthate as a collector.

[0071] Each fraction was then dried and weighed and analyzed for gold.The +{fraction (1/8)} to −{fraction (1/4)} inch material represented 51%of the weight and 18% of the gold at 0.48 ppm Au. The +30 mesh to−{fraction (1/8)} inch material represented 28% of the weight and 32% ofthe gold at 1.47 ppm Au. The total pyrite, which included both thegravity separated pyrite and the pyrite concentrate from the flotationof the sand, represented 4.7% of the weight and 35% of the gold at 9.8ppm Au. The remaining sand flotation tail and clay material represented16% of the weight and 14.6% of the gold at about 1.2 ppm Au.

[0072] The +{fraction (1/8)} to −{fraction (1/4)} inch material and the+30 mesh to −{fraction (1/8)} inch material were combined according totheir weight percentages. The combined material was adjusted to a pH of2.0 with 10% sulfuric acid at 30 ml/kg. The one mixture was then pouredinto a column and aerated from the bottom with at least 15 l ofair/min/m² and liquid dilute basal solutions of (NH₄)₂SO₄ 0.04 g/lMgSO₄.7H₂O at 0.04 g/l and KH₂PO₄ at 0.004 g/l were added to the top atabout 15 ml/hour. Thiobacillus ferrooxidans bacteria were added to thetop of the column and washed into the column with the liquid flow. Thisprocedure allowed for good air flow and liquid flow and also migrationof bacteria through the column. After about one month the effluent fromthe column showed good bioleaching of iron at about 0.1% per day.

EXAMPLE 2

[0073] A second sample of ore from the same mine as in Example 1 wascrushed to −{fraction (3/8)} inches. Four 23 Kg splits of this samplewere combined and wet screened into a +{fraction (1/4)} inch, a+{fraction (1/8)} to −{fraction (1/4)} inch, a +10 mesh to −{fraction(1/8)} inch, a +16 mesh to −10 mesh, a +30 to −16 mesh, a +60 to −30mesh, and a −60 mesh fraction. The +60 to −30 mesh and the −60 meshfraction were used to evaluate a number of gravity separations to make apyrite fraction a sand fraction and a clay fraction. The dry weights ofeach size fraction were used to calculate the weight percentage of thesize fraction. Each size fraction was also analyzed for the amount ofgold, iron and gold extraction by traditional cyanide leaching (seeTable 1).

[0074] The five size fractions larger than 30 mesh were put intoindividual columns for biooxidation. Bacteria and nutrients were addedas in Example 1 and air was blown in from the bottom or top of thecolumn. The progress of the biooxidation was monitored by measuring theamount of iron leached from the ore by using atomic absorption analysisof the nutrient solution passing through the column. The approximatetotal amount of iron in each size fraction of the ore was calculatedfrom the weight of the size fraction and an iron analysis of arepresentative sample of the ore. The percent iron leached and theaverage percent iron leached each day are plotted against time for allfive size fractions in FIGS. 2 and 3, respectively. TABLE 1 ORE SIZEFRACTION ANALYSIS GRAVITY SEPARATION Au Au % BIOOX. % SIZE WT % (wt %)(ppm) Fe % REC. RECOV. +¼ 20.9 0.57 2.4 24.3 50.6 (15) ⅛-¼ 32.3 0.78 2.638.8 62.7 (24) 10-⅛ 4.89 0.525 3.8 47.3 76.1 (40) 16-10 8.49 1.22 3.844.3 74.7 (46) 30-16 9.36 1.92 5.8 37.3 84.4 (53) 60-30 6.65 pyrite 1.6%13.56 47.1 sand 5.02% 0.43 75.3 -60 17.3 pyrite 2.68% 7.81 69.9 clay14.62% 1.48 86.5

[0075] After several months of biooxidation, samples were taken fromeach column and the percent iron leached noted. The partiallybiooxidized ore was then leached with cyanide in the same way theoriginal unoxidized samples were. The gold extraction of the unoxidizedsample and the biooxidized sample are compared in Table 1. The percentbiooxidation for each size fraction is reported in Table 1 inparentheses. From this data one can see that the smaller size fractionsbiooxidized at a faster rate. Also, all the size fractions show anincrease in gold extraction after being biooxidized.

[0076] The +60 to −30 mesh and −60 mesh size fractions were alsoanalyzed for gold extraction. The sand tails from a shaker tableseparation of the refractory pyrite from the +60 to −30 mesh fractionwas fairly low in gold, but the gold was cyanide extractable withoutbiooxidation (75%). The very fine sand and clay from the −60 meshfraction was higher in gold and in gold extraction (86%). This indicatedthat no further oxidation of the very fine sand and clay materials inthis size fraction was required.

[0077] The removal of the small size fractions (i.e., the size fractionshaving a particle size less than 30 mesh) including the clay fractionallowed all the columns to have excellent air flow. Columns made withwhole ore or whole ore with agglomeration often would become plugged,inhibiting air flow. Thus, by separating the fines and clays, largescale heaps may be constructed without having to use larger crush sizes(i.e., {fraction (3/4)} inch or larger) to achieve good air flow.

[0078] The pyrite fractions of the −30 and −60 mesh fractions were bothhigh in gold and refractory to cyanide leaching. These pyrite fractionswere combined and then milled to −200 mesh in a ball mill. The −200 meshpyrite concentrate was used in shake flask experiments to determine theamount of gold extraction as a function of percent biooxidation (seeFIG. 4). In preparing these tests, 75 ml of a 500 ppm cyanide solutionwas added to 30 gm of the pyrite concentrate. The solution and ore wasthen rolled at 10 rpm for 96 hrs. before the cyanide solution was testedto determine the amount of gold extracted.

[0079] Some of the pyrite from the gravity separated fines was furtherprocessed by grinding to −200 mesh and floating with xanthate to from aconcentrate of over 50% pyrite. A sample of this concentrate weighing500 gm was then mixed with 500 ml of solution containing iron oxidizingbacteria at greater than 108 cells per ml and 3000 ppm ferric sulfate.After one hour, the 500 gm sample of pyrite concentrate suspended in 500ml of ferric-bacteria solution was poured directly onto the top of the+{fraction (1/4)} inch ore column, containing about 15 Kg of ore. Thiswas done after biooxidation of the column ore had been in progress forover 300 days. The black liquid spread quickly down through the columnwith most of the pyrite concentrate being retained by the column. Thesmall amount of pyrite concentrate that did pass through the column waspoured back onto the top of the column and was retained by the column onthe second pass. The pyrite appeared to be evenly distributed throughoutthe column and did not inhibit the air flow.

[0080] Liquid at pH 1.8 was dripped onto the top of the column, as hadbeen done throughout the experiment. The flow rate was about 200 ml perday. The liquid collected after three days had dropped in Eh from about650 mV to 560 mV. The pH was still at about 1.8 as it had been for along time. The iron concentration in the liquid was 2800 ppm, which wasjust a little lower than the iron concentration of the added bacteriasolution. Two days after adding the pyrite concentrate to the column,the iron concentration in the off solution had increased to 4000 ppm andthe pH had dropped to 1.6 indicating that biooxidation of the pyrite hadstarted. FIGS. 5, 6, and 7 illustrate the change in Eh, pH, and ironconcentration of the column effluent, respectively over time.

[0081] Another aspect of the present invention will now be described. Inthis aspect, a process for recovering precious metal values from aconcentrate of precious metal bearing refractory sulfide minerals isdescribed. The process comprises (a.) distributing a concentrate ofrefractory sulfide minerals on top of a heap of support material; (b.)biooxidizing the concentrate of refractory sulfide minerals; (c.)leaching precious metal values from the biooxidized refractory sulfideminerals with a lixiviant; and (d.) recovering precious metal valuesfrom the lixiviant.

[0082] A concentrate of precious metal bearing refractory sulfideminerals will typically be prepared from a precious metal bearingrefractory sulfide ore. The concentrate can be prepared from such oresusing well known gravity separation or flotation techniques. Althoughgravity separation is cheaper, flotation is the preferred method ofseparation because of the selectivity of the process. The mostfrequently used collector for concentrating sulfide minerals in aflotation process is Xanthate. Xanthate flotation processes are wellknown to those skilled in the art and need not be described in detailherein.

[0083] Preferably the particle size of the concentrate is such that 80to 90% of the concentrate is less than 100 to 300 mesh. More preferably,80 to 90% of the concentrate is less than 100 to 150 mesh.

[0084] The optimum size may, however, vary with various ore types. Ingeneral, the operator should strive for a particle size which permitsoptimum separation in the concentration process and which provides forthe optimal rate of biooxidation versus the incremental costs ofadditional fine grinding.

[0085] The smaller the particle size of the sulfide minerals within theconcentrate, the more quickly the concentrate will oxidize duringbioleaching. However, the faster biooxidation rate does not alwaysjustify the added energy costs associated with fine grinding an ore or aflotation concentrate.

[0086] With the process according to the present invention, the cost ofleaving the concentrate on the heap to biooxidize is minimal. Therefore,a slightly longer biooxidation period may be justified to avoid havingto incur additional grinding related expenses. In this regard, thepresent process has an advantage over mill type processes. In mill typeprocesses, the sulfide mineral concentrate must be very finely ground toensure high biooxidation rates so that the bioreactor can process asmuch concentrate as possible in as short of period of time as possibleto maintain the economics of the process.

[0087] After the sulfide mineral concentrate is formed, it isdistributed over the top of a heap of support material. Preferably, theconcentrate is distributed on top of the heap in a slurry form so thatthe concentrate can be piped directly to the heap without having to bedried first. The pulp density of the concentrate should be adjusted sothat the concentrate flows well, but does not simply wash through theheap of support material. Because the sulfide mineral particles arehydrophobic, they will tend to stick to the support material rather thanmigrating completely through the heap if the appropriate supportmaterial is selected. Nor should blockage of flow channels be a problemif an appropriate size support material is selected.

[0088] The purpose of the support material is to capture and retain thesulfide minerals as they slowly migrate down through the heap so thatthe support material acts as a large surface area bioreactor. For thisreason support materials having a high degree of porosity or a roughsurface are preferred since these types of surfaces will tend to captureand retain the concentrate. The more concentrate that the support rockcan support without blockage of the flow channels the better. Supportmaterials that can be used in practicing the present invention includecoarse ore particles, lava rock, gravel, or rock containing smallamounts of mineral carbonate as a source of CO₂ for the biooxidizingbacteria. Lava rock is a particularly preferred support material due toits roughness and high degree of porosity.

[0089] Support material which contains a small amount of mineralcarbonate is beneficial not only for the CO₂ that it produces but isalso beneficial because it will help buffer the acid solution producedas a result of the biooxidation process. This will make it easier tocontrol the pH of the bioreactor during the biooxidation process.

[0090] With respect to selection of an appropriate size of supportmaterial, there are several competing interests that should beconsidered. Smaller diameter support materials have greater surface areaand thus increase the effective area of the bioreactor created by theheap of support material. However, smaller diameter support material maybe more expensive depending on the amount of grinding required toproduce the desired size. Further, smaller diameter support material maybe subject to more blockage of fluid flow channels by the concentratewhich is added to the top of the heap. Larger support material willpermit taller heaps to be formed without risk of flow channels becomingplugged.

[0091] Typically, the support material will be larger than about{fraction (1/4)} inch in diameter and smaller than about 1 inch indiameter. Preferably the support material is greater than about{fraction (3/8)} inch in diameter and less than about {fraction (3/4)}inch in diameter. A support material having a diameter of about{fraction (1/2)} inch should be the optimum size.

[0092] To biooxidize the concentrate, the heap is inoculated withbacteria or other microbe capable of biooxidizing the sulfide mineralsin the concentrate. Such microbial treatments are well known in the art.Bacteria that can be used for this purpose include Thiobacillusferrooxidans, Leptospirillum ferrooxidans, and Thiobacillus thiooxidans.Thiobacillus ferrooxidans is an especially preferred microorganism forbiooxidation processes.

[0093] If the bioleachate solution is recycled, precautionary steps maybe required to prevent toxic materials from building up in the recycledsolution so that the rate of biooxidation is not retarded significantly.The process described in U.S. patent application Ser. No. 08/329,002,filed Oct. 25, 1994, can be used to ensure that inhibitory materials donot build up to the point that they become detrimental to thebiooxidation process.

[0094] After the refractory sulfide concentrate is sufficientlybiooxidized, the liberated precious metal values can be leached with alixiviant of thiourea or cyanide. Cyanide is the preferred lixivianteven though the pH of the heap must first be raised prior to leaching.An advantage of thiourea is that it is not toxic to the biooxidizingmicroorganisms. As a result, intermittent leachings can be performed todissolve the liberated precious metal values and then the biooxidationprocess can be resumed.

[0095] Dissolved precious metal values can be recovered from thelixiviant using well known techniques to those skilled in the art suchas carbon in leach and carbon in column processes.

[0096] Another advantage of the present process is that it can be usedas a continuous process by intermittently adding fresh or newconcentrate to the top of the heap. The advantage of adding freshconcentrate to the top of the heap is that once the heap is establishedand biooxidation is occurring rapidly, the fresh concentrate can beadded to maintain the high rate of biooxidation within the heap withouthaving to tear down the heap to process the biooxidized material.

[0097] Due to the relatively low capital and operating costs of thepresent process, it can be used to economically process much lower gradeconcentrates, and as a result lower grade ores, than a mill biooxidationprocess. Further, by distributing the concentrate of precious metalbearing refractory sulfide minerals on top of a heap of supportmaterial, good fluid flow (both air and liquid) is ensured within theheap.

[0098] Another aspect of the present invention is now described. In thisaspect, a process is provided for recovering base metal values fromsulfide ores. Such ores include, by way of example, chalcopyrite,sphalorite, nickel sulfide ores, and uranium sulfide ores. The processaccording to this aspect of the invention comprises (a.) forming aconcentrate of metal sulfide minerals; (b.) spreading the concentrate ontop of a heap of support material; (c.) biooxidizing the concentrate;and (d.) recovering metal values from the solution used to biooxidizethe metal sulfide minerals. Due to the fact that this process, like theprocess previously described for processing concentrates of preciousmetal bearing sulfide minerals, uses a heap of support material for thebioreactor, its capital and operating costs are less than that of a millbioleaching operation. However, due to the good air flow in the heap,the biooxidation rate of the sulfide minerals is quite high and canapproach that of what would be observed in a mill type operation.

[0099] Depending on the sulfide ore from which the concentrate isobtained, the base metal values that can be recovered from the processaccording to the present aspect of the invention include copper, zinc,nickel and uranium.

[0100] The process parameters and considerations for the processaccording to the present aspect are much the same as those set forthabove for the method of processing precious metal bearing concentratesof refractory sulfide minerals. The primary difference between the twoprocesses, however, is that the base metal values of interest in thepresent process dissolve during the biooxidation process. As a result,the metal values are recoverable directly from the solution used tobiooxidize the concentrate of metal sulfide minerals. The technique usedto extract the metal values of interest will depend on the specificmetal of interest. As those skilled in the art will immediatelyrecognize, such techniques may include solvent extraction, ironcementation, and precipitation through pH adjustments. Solventextraction is a particularly preferred method of removing copper fromthe bioleachate solution.

[0101] As with the above described process for recovering precious metalvalues from a precious metal bearing concentrate of sulfide minerals,the present process can be operated in a continuous mode by addingconcentrate on an intermittent basis. For example, concentrate can beadded on a daily or weekly basis. As described above, such additionswill ensure that the rate of biooxidation remains high for theconcentrate that is distributed over the heap and which has migratedthrough the heap.

[0102] As one skilled in the art will recognize, the process accordingto the present aspect of the invention can be combined with the aboveprocess for recovering precious metal values from a concentrate ofrefractory sulfide minerals. This is because base metal values from therefractory sulfide minerals will inherently dissolve into thebioleachate solution during the biooxidation process whilesimultaneously liberating any occluded precious metal values in thesulfide minerals. These values can then be recovered if desired usingthe techniques described above.

EXAMPLE 3

[0103] Two simultaneous bioleaching tests were set up to test the rateof biooxidation of a gold bearing ore pyrite concentrate. The first testconsisted of a column type experiment to simulate a heap leachingprocess and the second consisted of a shake flask experiment to simulatea stirred tank process.

[0104] The starting concentrate for both tests was obtained from theJamestown mine in Tuolumne County, California. The mine is owned bySonora Gold corporation and lies along the mother lode vein system. Theconcentrate was produced using a xanthate flotation process andcontained 39.8% sulfides and 36.6% iron. The sulfide minerals within theconcentrate primarily consisted of pyrite. Size analysis showed thatover 76% of the concentrate particles were smaller than 200 mesh. Theconcentrate had a high gold concentration (about 2 oz. per ton ofconcentrate) and was known to be refractory to cyanide leaching.

[0105] The percentage of biooxidation in each of the tests wasdetermined by analysis of the iron concentration in all solutionsremoved from the column or in the case of the flask experiment theconcentration of iron in solution plus any iron solution removed.

[0106] A culture of Thiobacillus ferrooxidans was used to biooxidize thesulfide mineral concentrate in each of the tests. The culture ofThiobacillus ferrooxidans was originally started with ATCC strains 19859and 33020. The culture was grown in an acidic nutrient solution having apH of 1.7 to 1.9 and containing 5 g/l ammonium sulfate ((NH₄)₂SO₄)),0.833 g/l magnesium sulfate heptahydrate (MgSO₄.7H₂O), and 20 g/l ironin the form of ferrous and ferric sulfate. The pH of the solution wasadjusted to the above range using sulfuric acid (H₂SO₄).

[0107] Prior to application of the culture to the test samples, themixed culture of sulfide mineral oxidizing bacteria was grown to a celldensity of 4×10⁹ to 1×10¹⁰ cell per ml.

[0108] The column experiment was started by inoculating a 150 g sampleof concentrate with about 10⁸ cells per gram of concentrate. This wasaccomplished by adding three milliliters of bacteria at 5×10⁹ cells permilliliter to the 150 g sample of pyrite concentrate. The 150 g ofpyrite concentrate suspension was then poured into a 3 inch by 6 footcolumn filled about halfway with 3 liters of {fraction (3/8)} inch lavarock. The lava rock support material was chosen because it has a highsurface area and it holds up well to the acid condition encounteredduring biooxidation.

[0109] During inoculation and subsequent solution additions, the pyriteconcentrate did not wash out of the column. Most of the pyriteconcentrate was held in the first foot of the lava rock. Air and liquidwere introduced through the top of the column. The bioleach solution wasrecirculated until the pH of the column was adjusted down to about 1.8.After biooxidation started within the column, a 0.2×9K salts solutionhaving a pH of 1.8 and containing 2000 ppm of iron, primarily in theferric form, was fed to the column. The 2,000 ppm of iron was subtractedfrom all analysis of iron in solution coming off of the column.

[0110] The composition of the standard 9K salts medium for T.ferrooxidans is listed below. The concentrations are provided ingrams/liter. (NH₄)SO₄ 5 KCl 0.17 K₂HPO₄ 0.083 MgSO₄ · 7H₂O 0.833 Ca(NO₃)· 4H₂O 0.024

[0111] The notation 0.2×9K salts indicates that the 9K salt solutionstrength was at twenty percent that of the standard 9K salts medium.

[0112] After 26 days of biooxidation, about 35% of the iron in thepyrite concentrate had been oxidized. At this point, the test wasconverted to a continuous process test by adding 3 g of new concentrateto the column every day. After 9 more days, the rate of pyrite additionwas increased to 6 g per day.

[0113] The flask experiment was started at the same time as the columnexperiment. To start the experiment, a 50 g sample of the pyriteconcentrate was inoculated with 1 milliliter of the bacteria culture.The pyrite concentrate was then added to 1 liter of 0.2×9K saltssolution having a pH of 1.8 in a large shake flask. Not only was theconcentrate inoculated with the same bacteria, but it was alsoinoculated at the same number of cells per gram.

[0114] Air was introduced to the bioleach solution by orbital shaking ofthe flask at about 250 rpm. Solution was removed from the flask fromtime to time to keep the ferric concentration from getting much higherthan that in the column.

[0115] When the column experiment was converted to a continuous processon day 26, the flask experiment was also converted to a continuous testby adding 1 g of pyrite concentrate per day to the flask. After 9 moredays, the amount of concentrate added was increased to 2 g per day.

[0116] After 58 days, the pyrite additions to both the flask and columnexperiments were stopped. Both the column and the flask were thenallowed to biooxidize an additional 20 days. At this point, theconcentrate in the column was about 76% oxidized and the concentrate inthe flask was about 89% oxidized. The column was then leached for 10days with thiourea to extract liberated gold. The thiourea onlyextracted about 30% of the gold. However, after 3 days of reverting backto additions of the 0.2×9K salts solution having a pH of 1.8 andcontaining 2,000 ppm of ferric iron, the Eh and the iron concentrationof the column effluent increased. This indicated that the thiourea wasnot toxic to the bacteria and that thiourea extractions could be donefrom time to time without killing the bacteria.

[0117]FIG. 8 shows the amount of biooxidation versus time in days forboth the column and flask concentrate bioleaching tests. The phrase “TUleach” in FIG. 8 stands for thiourea leach. The data used to prepareFIG. 8 is tabulated in Tables 2 and 3 at the end of this example.

[0118] As indicated above, the flask was meant to simulate a stirredtank process. When the flask test was converted to a continuous processby adding pyrite each day, it was meant to simulate a large scaleprocess in which new pyrite is introduced on an intermittent basis to arapidly biooxidizing tank containing a large amount of bacteria thathave adapted to the ore. The daily addition of pyrite to the column wasdone to test the feasibility of a continuous process in which aconcentrate of precious metal bearing sulfide minerals is continuouslyor intermittently added to the top of a heap comprised of biooxidizingconcentrate distributed on a heap of support material such as lava rock.

[0119] As the above tests demonstrate, the rates of biooxidation werenot significantly different between the column and flask tests. Thestart of biooxidation was a little slower in the column test. This mayhave been due to about a 10 day lag time in adjusting the pH of thecolumn down to 1.8. The rate of biooxidation in the column then pickedup to be the same as the flask. Later in the experiment the rate beganto slow down again. This may have been due to a lack of mixing of thefresh pyrite with the biooxidizing pyrite. However, the rates ofbiooxidation between the two tests were close enough to demonstrate theviability of the process according to the present invention. Theviability of the present process is especially attractive in view of themuch lower capital and operating costs of a heap process as compared toa stirred tank process. TABLE 2 Data From Column Biooxidation Test %bioox. G. of iron Total g. of pyrite Conc. of Time in added to of ironbased on iron in days column removed iron g./l 0 54.400 2.830 5.2001.884 15 54.400 5.500 10.100 2.840 20 54.400 12.617 23.180 4.704 2354.400 14.480 26.620 4.976 26 55.540 19.230 34.620 9.088 27 56.63020.430 36.070 9.432 28 57.720 22.329 38.700 9.800 29 58.800 23.98740.800 6.400 30 59.900 25.176 42.000 5.964 31 61.000 27.075 44.380 5.87632 62.070 28.337 45.650 6.508 33 63.160 29.285 46.360 6.212 34 64.25030.257 47.080 4.900 35 65.340 31.824 48.700 7.224 36 69.700 32.97047.300 5.428 37 69.700 34.066 48.900 5.265 38 74.050 35.184 47.500 5.62039 74.050 36.302 49.000 40 76.230 37.420 49.100 5.120 41 78.410 38.42549.000 5.000 42 80.590 39.453 48.900 5.024 43 82.760 40.744 49.200 5.53644 84.940 42.172 49.600 5.808 45 89.300 43.602 48.800 5.964 46 89.30045.032 50.400 47 91.480 46.462 50.800 5.976 48 93.660 47.932 51.1806.200 49 95.836 49.650 51.800 6.896 50 98.014 50.582 51.600 7.328 51100.192 52.142 52.040 8.240 53 104.548 55.591 53.170 9.664 54 106.72658.012 54.360 8.052 55 108.896 59.835 64.950 8.288 56 111.066 61.57155.440 8.200 57 113.236 63.136 55.760 7.304 58 115.406 65.370 56.6408.384 59 115.406 67.640 58.610 8.484 61 115.406 70.806 61.350 8.208 62115.400 72.344 62.690 7.128 63 115.400 72.777 63.930 6.776 64 115.40075.013 65.000 5.852 65 115.400 76.169 66.000 5.728 66 115.406 77.32567.000 5.728 68 115.406 79.668 69.030 5.748 69 115.406 80.468 69.7304.668 70 115.400 81.043 70.220 4.740 71 115.400 81.828 70.904 4.856 72115.400 82.716 71.674 5.064 73 115.400 83.781 72.590 4.804 75 115.40084.975 73.630 4.488 76 115.400 85.609 74.180 4.112 90 115.400 87.17075.533 2.892 93 115.400 88.754 76.900 3.476

[0120] TABLE 3 Flask Biooxidation Data Flask % Time in bioox. by daysiron 0 4.930 15 18.890 21 29.850 26 37.400 28 39.790 36 45.370 44 46.89047 52.310 49 54.510 51 58.380 54 62.010 56 62.630 58 65.400 63 72.110 6981.410 72 83.300 75 89.440

[0121] Although the invention has been described with reference topreferred embodiments and specific examples, it will readily beappreciated by those of ordinary skill in the art that manymodifications and adaptions of the invention are possible withoutdeparture from the spirit and scope of the invention as claimedhereinafter. For example, while some of the processes according to thepresent invention have been described in terms of recovering gold fromrefractory sulfide or refractory carbonaceous sulfide ores, theprocesses are equally applicable to other precious metals found in theseores such as silver and platinum.

I claim:
 1. A process for recovering precious metal values from a concentrate of precious metal bearing refractory sulfide minerals, comprising: a. distributing the concentrate of refractory sulfide minerals on top of a heap of support material; b. biooxidizing the concentrate of refractory sulfide minerals; c. leaching precious metal values from the biooxidized refractory sulfide minerals with a lixiviant; and d. recovering precious metal values from the lixiviant.
 2. A process according to claim 1, wherein the precious metal recovered from the lixiviant is at least one selected from the group consisting of gold, silver and platinum.
 3. A process according to claim 1, wherein the precious metal recovered from the lixiviant is gold.
 4. A process according to claim 1, wherein the support material is selected from the group consisting of lava rock, gravel, and coarsely ground ore.
 5. A process according to claim 1, wherein the support material is lava rock.
 6. A process according to claim 1, wherein the lixiviant is selected from the group consisting of thiourea and cyanide.
 7. A process according to claim 1, wherein the lixiviant is thiourea.
 8. A process according to claim 1, further comprising adding fresh concentrate to the top of the heap on an intermittent basis.
 9. A process according to claim 8, wherein the precious metal values are intermittently leached from the biooxidized refractory sulfide minerals with thiourea.
 10. A process for recovering precious metal values from a concentrate of precious metal bearing refractory sulfide minerals, comprising: a. distributing the concentrate of refractory sulfide minerals on top of a heap of support material, wherein the support material is selected from the group consisting of lava rock, gravel, and coarsely ground ore; b. biooxidizing the concentrate of refractory sulfide minerals; c. leaching precious metal values from the biooxidized refractory sulfide minerals with a lixiviant; and d. recovering precious metal values from the lixiviant.
 11. A method according to claim 10, wherein the precious metal recovered is selected from the group consisting of gold, silver, and platinum.
 12. A method according to claim 10, wherein the lixiviant is selected from the group consisting of thiourea and cyanide.
 13. A process according to claim 10, further comprising adding fresh concentrate to the top of the heap on an intermittent basis.
 14. A process according to claim 10, wherein the precious metal values are intermittently leached from the biooxidized refractory sulfide minerals with thiourea.
 15. A process for recovering gold values from a concentrate of gold bearing refractory sulfide minerals, comprising: a. distributing the concentrate of refractory sulfide minerals on top of a heap of support material, wherein the support material is selected from the group consisting of lava rock, gravel, and coarsely ground ore; b. biooxidizing the concentrate of refractory sulfide minerals; c. adding fresh concentrate to the top of the heap on an intermittent basis; d. intermittently leaching gold from the biooxidized refractory sulfide minerals with thiourea; and e. recovering gold values from the thiourea.
 16. A process for recovering metal values from a sulfide ore, comprising: a. forming a concentrate of metal sulfide minerals; b. spreading the concentrate on top of a heap of support material; c. biooxidizing the concentrate; and d. recovering metal values from the solution used to biooxidize the metal sulfide minerals.
 17. A method according to claim 16, wherein the metal values recovered are selected from the group consisting of copper, zinc, nickel, and uranium.
 18. A method according to claim 16, wherein the metal recovered is copper.
 19. A process according to claim 16, wherein the support material is selected from the group consisting of lava rock, gravel, and coarsely ground rock.
 20. A process according to claim 16, wherein the support material is lava rock.
 21. A process according to claim 16, further comprising adding fresh concentrate to the top of the heap on an intermittent basis.
 22. A process for recovering metal values from a sulfide ore, comprising: a. forming a concentrate of metal sulfide minerals; b. spreading the concentrate on top of a heap of support material, wherein the support material is selected from the group consisting of lava rock, gravel, and coarsely ground rock; c. biooxidizing the concentrate; d. adding fresh concentrate to the top of the heap on an intermittent basis; and e. recovering metal values from the solution used to biooxidize the metal sulfide minerals.
 23. A process according to claim 22, wherein the support material is lava rock.
 24. A process according to claim 22, wherein the metal recovered is selected from the group consisting of copper, zinc, nickel, and uranium.
 25. A process according to claim 22, wherein the metal recovered is copper.
 26. A method for recovering metal values from refractory sulfide ores comprised of metal sulfide particles, the process comprising the steps of a. separating fines from a crushed refractory sulfide ore; b. forming a heap with said refractory sulfide ore; c. bioleaching the ore in said heap to thereby oxidize the metal sulfide particles contained therein; d. hydrometallurgically treating the bioleached ore to recover metal values; and f. treating the separated fines to recover metal values contained therein.
 27. A method for recovering precious metal values from refractory sulfide ores comprised of metal sulfide particles having occluded precious metal values, the process comprising the steps of: a. separating fines from a crushed refractory sulfide ore; b. forming a heap with said refractory sulfide ore; c. bioleaching the ore in said heap to thereby oxidize the metal sulfide particles contained therein; d. hydrometallurigically treating the bioleached ore to recover precious metal values; and e. treating the separated fines to recover precious metal values contained therein.
 28. A method according to claim 27, wherein said method of fines treatment comprises: a. separating precious metal containing metal sulfide particles from the fines; b. oxidizing said metal sulfide particles; and c. hydrometallurgically treating said oxidized metal sulfide particles to recover precious metal values contained therein.
 29. A method according to claim 28, further comprising: a. agglomerating the fines after separation of said metal sulfide particles; and b. hydrometallurgically treating said agglomerated fines to recover precious metal values.
 30. A method according to claim 27, wherein said method of fines treatment comprises: a. separating precious metal containing metal sulfide particles from the fines; and b. adding said metal sulfide particles to the heap.
 31. A method according to claim 30, further comprising: a. agglomerating the fines after separation of said metal sulfide particles; and b. hydrometallurgically treating said agglomerated fines to recover precious metal values.
 32. A met-hod according to claim 27, wherein said method of fines treatment comprises: a. separating precious metal containing metal sulfide particles from the fines; b. hydrometallurgically treating said metal sulfide particles to recover nonrefractory precious metal values; c. oxidizing said metal sulfide particles; and d. hydrometallurgically treating said oxidized metal sulfide particles to recover additional precious metal values.
 33. A method according to claim 32 further comprising: a. agglomerating the fines after separation of said metal sulfide particles; and b. hydrometallurgically treating said agglomerated fines to recover precious metal values.
 34. A method according to claim 27, wherein said method of fines treatment comprises: a. separating precious metal containing metal sulfide particles from the fines; b. hydrometallurgically treating said metal sulfide particles to recover nonrefractory precious metal values; and c. adding the hydrometallurgically treated metal sulfide particles to the heap.
 35. A method according to claim 34, further comprising: a. agglomerating the fines after separation of said metal sulfide particles; and b. hydrometallurgically treating said agglomerated fines to recover precious metal values.
 36. A method according to claim 27, wherein said hydrometallurgical treatment comprises leaching said heap with a lixiviant selected from the group consisting of cyanide and thiourea.
 37. A method according to claim 27, wherein said hydrometallurgical treatment comprises leaching said heap with cyanide.
 38. A method according to claim 27, wherein said crushed refractory sulfide ore has a maximum particle size in the range of approximately {fraction (1/4)} inch to 1 inch, and said fines have a maximum particle size of about −60 mesh to −{fraction (1/8)} inch.
 39. A method according to claim 27, wherein the recovered precious metal is at least one metal selected from the group consisting of gold, silver, and platinum.
 40. A method according to claim 27, wherein the recovered precious metal is gold.
 41. A method according to claim 28, wherein said separated metal sulfide particles arc oxidized by biooxidation.
 42. A method according to claim 32, wherein said separated metal sulfide particles are oxidized by biooxidation.
 43. A method according to claim 28, wherein said metal sulfide particles are separated from the fines by a method selected from the group consisting of gravity separation and flotation.
 44. A method according to claim 32, wherein said metal sulfide particles are separated from the fines by a method selected from the group consisting of gravity separation and flotation.
 45. A method according to claim 27, further comprising treating the bioleached ore to inhibit pregrobbing by carbonaceous components contained therein. 